Process for the preparation of potassium sulfate from kainite



A. SCARFI AL Filed May 17, 1967 H .m F N D 5.1 VIG W? W. x E .A/ E 1 Z w mm mm FIIIIQIIIilqIIIIL 2 \J Tlllllu B J a 2 O 2 h 5 7 3 5 7 rlP li 3 limHly 4 2 5 2 6 2 7 a B m l||| |..l|||l|| June 24, 1969 PROCESS FOR THE PREPARATION OF POTASSIUM SULFATE FROM KAINITE United States Patent Oflice 3,451,770 Patented June 24, 1969 PROCESS FOR THE PREPARATION OF POTASSIUM SULFATE FROM KAINITE Alberto Scarfi and Emanuele Gugliotta, Siracusa, Italy, assiguors to Montecatini Edison S.p.A., Milan, Italy Filed May 17, 1967, Ser. No. 639,130 Claims priority, application Italy, May 18, 1966, 11,344/ 66 Int. Cl. C01d /00 US. Cl. 23-121 6 Claims ABSTRACT OF THE DISCLOSURE Raw kainite is leached at temperatures greater than 90 C. with an epsomite brine into a langbeinite slurry, a portion which is thence reacted with a schoenite brine, whereby there results potassium chloride and epsomite which are separated from each other and an epsomite brine. A portion of the potassium chloride is reacted with magnesium sulfate in the presence of a sulfate brine thus affording schoenite and a schoenite brine which is recycled and the remaining potassium chloride is reacted with said schoenite in the presence of water, whereby there are obtained potassium sulfate and sulfate brine.

Background of the invention This invention relates to a process for the production of potassium sulfate, and, more particularly, this invention relates to a process for the production of potassium sulfate from raw kainite ore. It is known that raw kainite ore always contains a considerable amount of sodium chloride: the NaCl content of this ore generally ranges from to by weight. To avoid that the potassium sulfate and the other salts obtainable from kainite be contaminated by sodium salts, it is therefore necessary to eliminate most of the sodium chloride contained in the kainite before or during processing of the same.

The technique most commonly used for the above purpose consists in the flotation of ground kainite ore: kainite undergoes flotation under use of suitable collecting agents, while most of the sodium chloride remains in the flotation wastes. The kainite beneficiated through flotation generally still contains about 5% by weight of NaCl.

This flotation operation is very expensive but so far it is deemed unavoidable, as other processes of equal technical validity were not known for separating the sodium chloride.

Our Italian Patent No. 672,661 describes a process for obtaining salts of potassium and magnesium from kainite. According to said process, kainite is leached, at temperatures higher than 90 C., with brines having suitable composition, the result being the obtaining of langbeinite. This langbeinite is cooled in the presence of its mother liquor, to temperatures lower than 400 0; according to the final concentration of the brine by magnesium chloride and according to saturation conditions by potassium chloride and magnesium sulfate said cooling permits the formation of: (l) schoenite (or leonite) and potassium chloride, or (2) schoenite (or leonite) and MgSO -7H O, or (3) potassium chloride and MgSO -6H O and/or MgSO -7H O. This process permits one to obtain best yields of potassium, as the brines in equilibrium with the salt mixtures obtained through the cooling action have a low content of potassium and this is due to the high magnesium content of said brines. This high MgCl content has, moreover, another important consequence; it greatly limits the solubility of sodium chloride in brines. The quantity of sodium chloride which can be eliminated with such a process from the operational cycle through the drained brines is consequently very limited, therefore, also the quantity of NaCl which can be introduced into the operational cycle with the kainite is very limited, otherwise potassium and magnesium salts are contaminated with sodium salts. In practice, the kainite entering the cycle cannot contain more than 2.5% of sodium chloride, when langbeinite is converted to KCl and MgSO -6H O and/or MgSO -7H O while the same kainite cannot have more than 5% sodium chloride when langbeinite is converted to schoenite and KCl or schoenite and MgSO -7H O.

The result is, therefore, that the technique of preliminary flotation cannot itself afford a beneficiation sufiicient to permit the processing of kainite according to the high potassium yield process described in Italian Patent No. 672,661, when a conversation of langbeinite to KCl and MgSO 6H O and/or MgSO -7H O is: to be obtained. When, on the contrary, langbeinite is, according to the same patent, to be converted to schoenite and KCl or epsomite, the top beneficiation of the kainite obtainable through preliminary flotation strictly corresponds to the process requirements, which impose a very careful flotation. Said patent already describes an alternative to the preliminary flotation of kainite; the langbeinite obtained through hot leaching of raw kainite can be converted by cold process to a mixture of KCl, NaCl,

MgSO 6H O and/ or MgSO 7H O.

The magnesium sulfate can be separated from the chlorides by classification or screening, while the potassium chloride can be separated from the sodium chloride through flotation. This latter operation permits one to obtain potassium chloride containing less than 5% by weight of NaCl.

This flotation carried out within the cycle, is, however, rather expensive due, among other things, to the large quantity of actual sodium chloride; the weight ratio between KCl and NaCl is in fact generally about 0.5. To avoid the contamination of magnesium sulfate due to NaCl it is, moreover necessary to have the ore ground to a degree of fineness higher than that required for a hot leaching operation. The sodium chloride present in the kainite ore submitted to grinding suflicient for a hot leaching operation contains in fact a significant percent age off relatively large cyrstals. As the granulometry of sodium chloride practically undergoes no modification through the various process stages, the larger crystals are found in the classified magnesium sulfate. It is therefore necessary to care for a more effective grinding of the ore, such as to reduce the granulometry of the sodium chloride; this, consequently, involves a more expensive grinding operation.

One object of this invention is to obtain potassium sulfate from kainite employing a process which aifords very high yields of potassium. Another object of this invention is to obtain potassium sulfate from kainite only, without having to resort to any external source or potassium chloride.

A further object of this invention is to obtain potassium sulfate from raw kainite without having to resort to any preliminary beneficiation of the ore obtained through flotation.

A still further object of this invention is to eliminate most of the sodium chloride through a very simple and inexpensive physical operation carried out during the process cycle.

These and still other objects are achieved through the process of this invention and according to which it is possible to obtain potassium sulfate from raw kainite through the following process stages:

(l) The ground raw material is subjected to leaching, at a temperature higher than 90 C., with recycling epsomite brine obtained at stage 3; consequently, one obtains a langbeinite slurry consisting of a solid phase containing langbeinite and sodium chloride, and a langbeinite brine.

(2) The langbeinite slurry is subjected to classification through which most of the sodium chloride is separated from said slurry.

(3) The langbeinite slurry is afterwards cooled to a temperature ranging from 35 C. to 20 C., in the presence of a schoenite brine originating from stage 6, consequently forming a solid phase consisting of a mixture, including KCl, NaCl, MgSO .6H O and/or MgSO .7H O, and an epsomite brine.

(4) The magnesium sulfate is separated from the chlorides through classification or screening while the epsomite brine is partially conveyed to waste and partially recycled to stage 1. The magnesium sulfate is partially conveyed to stage 6 and partially removed from the cycle.

(5) The potassium chloride is separated through flotation from the sodium chloride.

(6) Part of the potassium chloride, obtained at stage 5, is permitted to react at a temperature comprised between 20 C. and 35 C. and in the presence of a sulfate brine coming from stage 7, with a portion of the magnesium sulfate obtained at stage 4; schoenite (or leonite) and a schoenite brine are thus obtained. Said schoenite brine is recycled to stage 3.

(7) The schoenite is permitted to react, at a temperature comprised between 20 C. and 40 C. and in the presence of water, with the remaining potassium chloride obtained at stage 5; potassium sulfate and sulfate brine are thus obtained. Said sulfate brine is recycled to stage 6.

The sodium chloride is, according to this invention, driven away from the cycle not only together with the drained epsomite brine, but also through two operations carried out at two distinct stages: the classification of the langbeinite slurry obtained via hot process and the flotation of the KCl obtained via cold process. It has been in fact found that the langbeinite obtained by hot leaching appears in the form of small crystals which can be easily separated from most of the sodium chloride through a simple classifying operation.

It is in this way possible to eliminate about from 70% to 80% of the sodium chloride present in the langbeinite slurry. The small crystals of sodium chloride remain in the langbeinite slurry. They can be therefore found again in the mixture of obtained salts by the cold process; during the cold classification or screening said crystals pass through together with potassium chloride and a small quantity of magnesium sulfate. A subsequent flotation operation permits one to easily separate the potassium chloride from the residual sodium chloride and magnesium sulfate, which are driven away from the cycle. This invention permits one to obtain a very high yield of potassium sulfate; said yield, calculated according to the ratio between K 0 leaving the cycle in the form of K SO and K 0 entering the cycle in the form of kainite, generally ranges from 84% to 85%. This yield is appreciably higher than the yield afforded by most common processes for the production of potassium sulfate from kainite, i.e., the process of a linked conversion and the process of a conversion followed by metathesis. The linked conversion process, which is based on the conversion of kainite to schoenite and the following conversion of the schoenite to K 80 and which does not require any external bringing in of KCl, gives a potassium yield of about 60%, calculated according to the ratio between the K 0 leaving the cycle in the form of K 80 and the K 0 coming into cycle as kainite. The conversion and metathesis process, which is based on the conversion of kainite into schoenite and following metathesis between schoenite and KCl introduced to the cycle, gives a potassium yield generally ranging from to calculated according to the ratio between the K 0 leaving the cycle in the form of K 50 and the K 0 coming into the cycle as kainite and KCl.

According to this invention only a small fraction of the potassium (about 11% by weight) is lost as a consequence of the drainage of a part of the epsomite brine. Also, the loss of potassium due to the solid phases driven away from the cycle (hot classification wastes, classification or screening wastes, and cold flotation wastes) is very limited. Among the three possible processes of the langbeinite slurry obtained via hot process conversion to schoenite and KCl, conversion to schoenite and MgSO .7H O and conversion to KCl and MgSO .6H O and/ or the process chosen by this invention, viz, the conversion of langbeinite to KCl and MgSO .6H O and/ or MgSO .7H O, leads to the formation of a mother liquor with a higher MgCl content and, consequently, with a lower potassium content, so this process is the most advantageous as to the potassium yield.

The brines in equilibrium with a solid phase consisting of KCl and MgSO .6H O and/or MgSO .7H O have in fact a content of magnesium chloride (ranging from 52 to 65 moles per 1000 moles water) clearly higher than the already high content (ranging from 45 to 52 moles per 1000 moles water) of the brine in equilibrium with a solid phase consisting of schoenite and KCl or schoenite and MgSO .7H O.

All the potassium chloride obtained by the cold conversion of the langbeinite and a part of the magnesium sulfate obtained at the same stage are afterwards reacted in two distinct stages, with intermediate formation of schoenite, to obtain potassium sulfate. With the process of this invention it is therefore possible to obtain potassium sulfate from kainite only, that is, no external source of KCl is required for carrying out the metathesis reaction between schoenite and KCl. The various stages of the process being the object of this invention are hereinunder described in better detail with reference also to the figure of drawing diagrammatically illustrating the proposed process.

The continuous lines in said figure refer to the solid phases, while the dashed lines refer to the brines.

First stageleaching of the kainite The raw kainite ore is crushed to pieces dimensioned less than the values between 1 and 10 mm. The ore is preferably crushed to pieces dimensioned less than the values between 5 and 8 mm.

Crushed ore 11 is led to leaching stage 1 together with a part 12 of epsomite brine 1718 obtained at stage 3. The leaching temperature is higher than C., preferably about C.

Second stageclassification of the langbeinite Langbeinite slurry 13 obtained at stage 1 and consisting of langbeinite, sodium chloride and langbeinite slurry, is conveyed to classification stage 2. This operation is carried out in such a way as to separate the solid substances with dimensions higher than the values included between 35 and 60 mesh (the mesh sizes are always referred to the Tyler scale) and preferably higher than the values included between 35 and 48 mesh. The coarse fraction 14 containing most of the sodium chloride, is driven away from cycle.

The fine fraction 15 contains the remainder of the langbeinite slurry, viz, langbeinite, the fine part of the sodium chloride and the langbeinite brine.

Third stageconversion of the langbeinite to potassium chloride and epsomite The after grading langbeinite slurry 15 is collected, to-

gether with schoenite brine 16, and cooled at stage 3 to a temperature comprised between 35 C. and 20 C. This cooling is preferably carried out by subsequent stages in order to achieve the best conditions for heat recovery.

A mixture of KCl, NaCl and MgSO -6H O and/or MgSO -7H O is obtained in this manner. If the temperature is close to the lower limit of the above said range, i.e. 20 C., the magnesium sulfate tends to precipitate in the form of epsomite; if the temperature is close to the upper limit, MgSO -6H O tends to precipitate. It is preferable to process at 20 C. in such way that the magnesium sulfate precipitates in the form of epsomite.

Fourth stage--screening or cold classification Slurry 17, obtained at stage 3, undergoes in stage 4 a screening or classifying process. This screening is commonly carried out by means of screens with 28 to 60 mesh but preferably 28 to 35 mesh. Top screening 20, consisting of MgSO -6H O and/or MgSO -7H O is divided in two parts; first part 21 is driven away from cycle, while second part 22 is conveyed to stage 6. The chlorides are separated from epsomite brine 18 which is divided in two parts; first part 12 is recycled to stage 1 while second part 19 is drained after having been submitted to further operation for the recovery of the potassium contained therein, if that is the case. 35% to 40% of the epsomite brine is preferably conveyed to waste.

Fifth stage-flotation of the potassium chloride Mixture 23 of potassium chloride and sodium chloride is conveyed to flotation stage 5. Fatty amines are preferably used as flotation agents. Flotation wastes 24, mainly consisting of sodium chloride, are conveyed to waste.

Sixth stage-production of schoenite (or leonite) A part of KCl 25 submitted to flotation is conveyed to the schoenite (or leonite) production stage 6 together with a part 22 of the epsomite (or MgSO -6H O).

The brine fed to this stage is a sulfate brine 26 coming from stage 7. The reaction between KCl and epsomite is carried out at a temperature comprised between 20 C. and 35 C., preferably at about 20 C. At the end of the reaction one obtains a mixture 27 consisting of schoenite (or leonite) and eventually KCl, which is conveyed to metathesis stage 7, and schoenite brine 16 which is recycled to stage 3.

Seventh stage--metathesis Mixture 27 of schoenite and eventually KCl is reacted with the residual KCl in the presence of water 28, at a temperature comprised between 20 C. and 40 C. but preferably about 30 C.

Instead of conveying the potassium chloride obtained at stage 5, partially to stage 6 and partially to stage 7, it is preferable to convey it Wholly to stage 6, as the enclosed figure shows. Potassium sulfate 29 and sulfate brine 26 are obtained after the metathesis operation, said brine being recycled to stage 6.

In order to control the composition of the brines used in the various operational stages in such manner as to bring them to a degree of concentration deemed the best for every stage, it is possible to modify the content of the brine entering every stage by adding water in such a quantity as to establish a brine with a preestablished optimum content. Said additions are not absolutely necessary, but allow one to steadily maintain the best process conditions as to the kinetics and the yield of every operation.

These additions allow one to easily adjust also the process conditions to any variation in kainite content of the incoming ore; the concentrations of the various brines can be thus maintained at optimum values in spite of the variations occurring in the characteristics of the incoming ore.

It is readily understood that the use of different techniques for the hot separating of the coarse fraction, as for example :by screening, cycloning or centrifugation of the sodium chloride from the langbeinite, permits one to obtain equally good results, provided that said techniques are suitably applied.

In order to further illustrate the present invention and the advantages thereof the following specific examples are given, it being understood that the same are merely intended to be illustrative and not limitative.

EXAMPLE 1 1) Leaching of the kainite ore at C.100 gr. kainite ore were crushed to 6 mm. pieces and consisting of 67.15 gr. kainite and 32.85 gr. NaCl, and these were treated for one hour contact time at 100 C. with 87.30 gr. of recycled brine coming from stage 3.

40.43 g r. langbeinite, 29.91 gr. NaCl, 201 gr. unconverted kainite and 114.95 gr. langbeinite brine were obtained.

(2) Classification of the langbeinite slurry at 100 C. l87.30 gr. langbeinite slurry obtained from stage 1 were graded at 35 mesh in a classifier.

25.94 gr. classification wastes were obtained as coarse fraction of the classification after centrifugation and washing, said wastes consisting of 2.01 gr. kainite and 23.93 gr. NaCl while 161.36 gr. langbeinite slurry, consisting of 40.43 gr. langbeinite, 5.98 gr. NaCl, 114.95 gr. langbeinite brine, were obtained as fine fraction.

(3) Cooling and decomposition of the langbeinite slurry at 20 C.-161.36 gr. langbeinite slurry, having the same composition as specified at point 2, were cooled to 20 C. with contemporaneous addition of 76.98 gr. mother liquor, coming from stage 6, and 5.61 gr. water. The obtained slurry was conditioned for four hours at 20 C.

22.16 gr. KCl, 69.49 gr. MgSO -7H O, 12.42 gr. NaCl and 139.88 gr. epsomite mother liquors were obtained. The potassium content of this mother liquor was 1.89% by weight, the sodium content amounted to 0.90%.

(45) Screening and flotation at 20 C.The slurry obtained at the previous stage (243.95 gr.) was screened at 28 mesh for eliminating as top screening most of the epsomite which, after centrifugation and washing, turned out to amount to 55.59 gr.

The bottom screening consisted of 22.16 gr. KCl, 12.42 gr. NaCl, 13.9 gr. MgSO -7H O and 139.88 gr. brine.

87.30 gr. brine were recycled to stage 1, while the remaining part (52.58 gr.) was driven away from cycle.

26.20 gr. concentrate (21.50 gr. KCl, 4.70 gr. NaCl) and 22.28 gr. wastes (13.90 gr. MgSO -7H O, 7.72 gr. NaCl, 0.66 gr. KCl) were obtained by flotation of the bottom screening. Said wastes were conveyed to drainage.

(6) Production of schoenite at 20 C.26.20 gr. sylvite concentrate and 39.94 gr. MgSO -7H O, drawn from the top screening, were reacted for a contact time of one and one half hours at 20 C. with 55.18 gr. sulfate mother liquors coming from stage 7.

44.34 gr. solid substances (29.26 gr. schoenite, 15.08 gr. NaCl) and 76.98 gr. schoenite mother liquors were obtained after centrifugation and washing.

The obtained schoenite and KCl were conveyed to stage 7, the brine to stage 3.

(7) Metathesis at 30 C.44.34 gr. solid substances, coming from stage 6, were treated for three hours at 30 C. with 33.43 gr. water. 22.59 gr. humid K 50 with composition K 40.21, Mg 0.89, CI 0.96, 80., 51.55, H O 6.38 by weight percent, and 55 .18 gr. sulfate mother liquor were obtained after centrifugation. The sulfate mother liquor was recycled to stage 6.

21.87 gr. K SO with K 0 content equal to 50.04% by weight were obtained after drying. The total process yield, calculated according to the ratio between K20 leaving the cycle in the form of K SO and K 0 entering the cycle in the form of kainite, was 84.50%.

7 EXAMPLE 2 (1) Leaching of the kainite at 100 C.-100 gr. kainite ore, crushed to 8 mm. pieces and consisting of 67.16 gr. kainite and 32.84 gr. NaCl, were treated for one hour contact time at 100 C. with 87.30 gr. recycled brine coming from stage 3. 39.95 gr. langbeinite, 2.01 gr. unconverted kainite, 30.33 gr. NaCl and 115.01 gr. langbeinite brine were obtained.

(2) Classification of the langbeinite slurry at 100 C.187.30 gr. langbeinite slurry obtained from stage 1 were graded at 48 mesh in a classifier. 27.79 gr. wastes (2.01 gr. kainite, 25.78 gr. NaCl) were obtained as coarse fraction of the classification after centrifugation and washing, while 159.51 gr. classified langbeinite slurry consisting of 39.95 gr. langbeinite, 4.55 gr. NaCl and 115.01 gr. langbeinite brine, were obtained as fine fraction.

(3) Cooling and decomposition of the langbeinite slurry at 20 C.l59.51 gr. langbeinite slurry having a composition the same as specified at point 2, were cooled to 20 C. under contemporaneous addition of 78.05 gr. schoenite mother liquors, coming from stage 6, and 5.48 gr. water.

The obtained slurry was conditioned for four hours at 20 C. 22.31 gr. KCl, 10.59 gr. NaCl, 69.92 gr.

and 140.21 gr. epsomite mother liquors were obtained. The potassium content of these mother liquors was 1.87% by weight, their sodium content amounted to 1.07%.

(45) Screening and flotation-The slurry obtained at the previous stage (243.04 gr.) was screened at 135 mesh for eliminating as top screening most of the epsomite which, after centrifugation and washing, turned out to amount to 48.94 gr. The bottom screening consisted of 22.31 gr. KCl, 10.59 gr. NaCl, 20.98gr. MgSO -7H O and 140.21 gr. epsomite mother liquors. 87.30 gr. mother liquors were recycled to stage 1 while the remaining part (52.91 gr.) was driven away from cycle. 26.29 gr. sylvite concentrate (21.72 gr. KCl and 4.97 gr. NaCl) and 27.19 gr. wastes (5.62 gr. NaCl, 0.59 gr. KCl, 20.98 gr. MgSO -7H O) were obtained by flotation of the bottom screening. Said wastes were conveyed to drainage.

(6) Production of schoenite at 20 C.26.69 gr. sylvite concentrate and 40.26 gr. MgSO -7H O, drawn from the top screening, were reacted for one and onehalf hours at 20 C. with 55.67 gr. sulfate mother liquors coming from stage 7. 29.41 gr. schoenite, 15.16 gr. KCl and 78.05 gr. schoenite mother liquors were obtained after centrifugation and washing. The obtained schoenite and KCl were conveyed to stage 6, the brine to stage 3.

(7) Metathesis at 30 C.44.57 gr. solid substance coming from stage 6, were treated for a contact time of three hours at 30 C. with 33.61 gr. water. 22.52 gr. humid K 50 with composition K 40.47, Mg 0.87, Cl 0.90, S0 51.94, H O 5.82 by weight percent, and 55.67 gr. sulfate mother liquors were obtained after centrifugation. The sulfate mother liquor was recycled to stage 6. 21.86 gr. K 50 with K 0 content equal to 50.22% by weight, were obtained after drying. The total process yield, calculated according to the ratio between K 0 leaving the cycle in the form of K 50 and K 0 entering the cycle as kainite, was 84.84%.

EXAMPLE 3 1) Leaching of the kainite ore at 100 C.-100 gr. kainite ore crushed to 66 mm. pieces and consisting of 71.70 gr. kainite, 28.30 gr. NaCl, were treated for one hour contact time at 100 C. with 87.30 gr. recycled brine coming from stage 3.

42.50 gr. langbeinite, 2.15 gr. unconverted kainite, 25.77 gr. NaCl and 116.88 gr. langbeinite brine were obtained.

(2) Classification of the langbeinite slurry at 100 C.187.30 gr. langbeinite slurry obtained from stage 1 were classified at 35 mesh in a classifier. 22.77 gr. wastes, consisting of 2.15 gr. kainite and 20.62 gr. NaCl, were obtained as coarse fraction of the classification after centrifugation and washing, while 164.53 gr. langbeinite slurry, consisting of 42.50 gr. langbeinite, 5.15 gr. NaCl 116.88 gr. langbeinite brine, were obtained as fine fraction.

(3) Cooling and decomposition of the langbeinite slurry at 20 C. 164.53 gr. langbeinite slurry having a composition the same as specified at point 2, were cooled to 20 C. under contemporaneous addition of 87.60 gr. schoenite mother liquors, coming from stage 5, and 5.92 gr. water. The obtained slurry was conditioned for 4 hours at 20 C. 24.93 gr. KCl, 11.58 gr. NaCl, 77.04 gr. MgSO -7H O, and 144.50 gr. epsomite mother liquor were obtained. The potassium content of this mother liquor was 1.91% by weight, the sodium content amounted to 1.11%.

(45) Screening and flotation-The slurry obtained at the previous stage (258.05 gr.) was screened at 28 mesh for eliminating as top screening most of the epsomite which, after centrifugation and washing turned out to amount to 61.63 gr.

The bottom screening consisted of 24.93 gr. KCl, 11.5 8 gr. NaCl, 15.1 gr. 'MgSO and 144.50 gr. epsomite mother liquor. 87.30 gr. mother liquor were recycled to stage 1 while the remaining part (57.20 gr.) was driven away from cycle.

29.93 gr. sylvite concentrate (24.43 gr. KCl, 5.50 gr. NaCl) and 21.99 gr. wastes (0.50 gr. KCl, 6.08 gr. NaCl, 15.41 gr. MgSO -7H O) were obtained by flotation of the bottom screening. Said wastes were conveyed to drainage.

(6) Production of schoenite at 30 C.-29.93 gr. sylvite concentrate and 45.37 gr. MgSO -7H O, drawn from the top screening, were reacted for one and one-half hours at 30 C. with 59.98 gr. sulfate mother liquor coming from stage 7. 31.28 gr. schoenite, 16.39 gr. KCl, and 87.60 gr. schoenite mother liquors were obtained after centrifugation and washing. Schoenite and KCl were conveyed to stage 7, the brine to stage 3.

(7) Metathesis at 30 C.47.67 gr. solid substances coming from stage 6 were treated for a contact time of three hours at 30 C. with 36.34 gr. water. 24.04 gr. humid KgSO with composition K 40.66, Mg 0.76, Cl 1.03, 80. 51.62, H O 5.92 by weight percent, and 59.98 gr. sulfate mother liquors were obtained after centrifugation.

The sulfate mother liquor was recycled to stage 6. 23.33 gr. K with K 0 content equal to 50.48% by weight, were obtained after drying.

The total process yield, calculated according to the ratio between K 0 leaving the cycle in the form of K 50 and K 0 entering the cycle as kainite, was 85.25%.

As many apparently widely different embodiments of the present invention can be made without departing from the spirit and scope thereof, it is to 'be understood that the same is not intended to be limited to the specific embodiments thereof, except as defined by the appended claims.

What is claimed is:

1. A process for the production of potassium sulfate from crushed raw kainite, comprising:

(a) leaching said crushed raw kainite at temperatures greater than about C. by means of epsomite brine recycled from below stage (c) thus affording a langbeinite slurry comprising a solid phase containing langbeinite and sodium chloride, and langbeinite brine;

(b) subjecting the said langbeinite slurry to classification and separating the bulk of the sodium chloride therefrom;

(c) cooling the said classified langbeinite slurry to a temperature of from between about 35 C. and about 20 C. in the presence of a schoenite brine recycled from below stage (f) thus affording a solid phase comprising a mixture of KCl, NaCl and a member selected from the group consisting of MgSO -6H O, MgSO -7H O and mixtures thereof, and epsomite brine;

(d) separating magnesium sulfate from chloride, partially draining the epsomite brine and partially recycling the same to above stage (a), partially conveying said magnesium sulfate to below stage (t) and partially expelling the same from the process cycle;

(e) separating potassium chloride from sodium chloride via flotation;

(f) reacting a portion of the potassium chloride obtained at stage (e) at a temperature from between about 20 C. and about 35 C. with magnesium sulfate from stage (d) in the presence of sulfate brine recycled from stage (g) thus affording a member selected from the group consisting of schoenite and leonite, and schoenite brine which is recycled to stage (c); and

(g) reacting the thus obtained stage (f) schoenite with the remaining potassium chloride from stage (e) at a temperature from between about 20 C. and about 40 C. in the presence of water thus affording the desired potassium sulfate and a sulfate brine which is recycled to stage (f).

2. The process as defined by claim 1, wherein stage ((1) the magnesium sulfate is separated by classification.

3. The process as defined by claim 1, wherein stage (d) the magnesium sulfate is separated by screening.

4. The process as defined by claim 1, wherein the raw kainite ore has been crushed to pieces of a size distribution of less than aobut 5 to 8 mm.

5. The process as defined by claim 1, wherein stage (b) langbeinite is classified such as to separate solids having a size distribution of greater than about 35 to 48 mesh.

6. The process as defined by claim .3, wherein stage (d) the member selected from the group consisting of MgSO -7H O and mixtures thereof, is separated from potassium chloride and sodium chloride by screening at a value ranging from between 28 to 35 mesh.

References Cited OSCAR R. VERTIZ, Primary Examiner. G. O. PETERS, Assistant Examiner.

UNITED STATES PATENT OFFICE Patent No.

Dated Inventor(s) June 24, 1969 Alberto Scarfi and Emaneule Gugliotta It is certified that error appears in the above-identified patent and that said Letters Patent are hereby corrected as shown below:

In the heading,

Column 1, line 57,

Column 2, line 46,

Column 2, line 46,

Column 7, line 32,

Column 10, line 7,

(SEAL) flan:

Edward M. Fletcher, It. Attesting Officer change change change change change change "11,344/66" to 18,034/66 L R EALET new low WILLIAM E. 'SGHUYIER, JER- oonmissioner of Patentu 

